Chloride assisted hydrometallurgical extraction of metal

ABSTRACT

A process for the extraction of a metal from an ore or concentrate comprises subjecting the ore or concentrate to pressure oxidation in the presence of oxygen and an acidic solution containing halogen ions and a source of bisulphate or sulphate ions, such as H2SO4. The metals which can be extracted by the process comprises copper as well as non-cuprous metals such as zinc, nickel and cobalt. During pressure oxidation the metal may be precipitated as an insoluble basic salt, such as basic copper sulphate, or substantially completely solubilized and precipitated later as the basic metal salt.

CROSS REFERENCE TO RELATED APPLICATION

This application is a continuation-in-part of U.S. patent applicationSer. No. 08/425,117 filed Apr. 21, 1995, now U.S. Pat. No. 5,645,708,the entire contents of which is incorporated herein by reference whichis a continuation-in-part of Ser. No. 08/098,874 filed Jul. 29, 1993 nowU.S. Pat. No. 5,431,788.

FIELD OF THE INVENTION

This invention relates to the hydrometallurgical treatment of metal oresor concentrates. In particular, it relates to the extraction of metalsfrom ores in the presence of halogen ions, such as chloride ions. Italso relates to the extraction of nickel and cobalt from laterite ores.

BACKGROUND OF THE INVENTION

Hydrometallurgical treatment of copper sulphide ores, such aschalcopyrite (CuFeS₂), is problematical because the severe conditionsrequired in a pressure oxidation step for the effective leaching ofcopper from these ores results in oxidation of the sulphide in the oreto sulphate, resulting in the generation of large amounts of acid whichrequires expensive neutralization. Attempts have been made to render thesulphide concentrate leachable under relatively milder conditions underwhich the sulphide would only be oxidized to elemental sulphur and notall the way through to sulphate. These attempts include the pretreatmentof the concentrate prior to the pressure oxidation step to render thesulphide concentrate more readily leachable, and the leaching of theconcentrate in the presence of chloride ions, such as described in U.S.Pat. No. 4,039,406. In this process, the copper values in theconcentrate are transformed into a solid basic copper sulphate fromwhich the copper values must then be subsequently recovered, asdescribed in U.S. Pat. No. 4,338,168. In the process described in U.S.Pat. No. 4,039,406 a significant amount (20-30%) of sulphide in the oreor concentrate is still oxidized to sulphate, resulting in greateroxygen demand during the pressure leach and the generation of sulphuricacid. This is particularly unfavourable for low grade concentrates,where the S/Cu ratio is high.

The present invention provides a process for the hydrometallurgicalextraction of copper and other metals in the presence of halogen ions,such as chloride and bromide in solution.

SUMMARY OF THE INVENTION

According to the invention, there is provided a process for theextraction of metal from a sulphide ore or concentrate, comprising thesteps of subjecting the ore or concentrate to pressure oxidation in thepresence of oxygen and an acidic chloride solution to obtain a resultingpressure oxidation filtrate and an insoluble basic metal sulphate salt,characterized in that the pressure oxidation is conducted in thepresence of a source of bisulphate or sulphate ions which is selectedfrom the group consisting of sulphuric acid and a metal sulphate whichhydrolyzes in the acidic solution and wherein the amount of the sourceof bisulphate or sulphate ions which is added contains at least thestoichiometric amount of sulphate or bisulphate ions required to producethe basic metal sulphate salt less the amount of sulphate generated insitu in the pressure oxidation.

According to one particular embodiment of the invention, the processfurther comprises the steps of recycling the pressure oxidation filtrateto the pressure oxidation step; leaching the basic metal sulphate saltproduced by the pressure oxidation in a second leaching with an acidicsulphate solution to dissolve the basic metal salt to produce a leachliquor containing metal sulphate in solution and a resulting solidresidue; separating the leach liquor from the solid residue; subjectingthe leach liquor to a solvent extraction process to produce metalconcentrate solution and a raffinate; and recycling the raffinate to thesecond leaching step. In this embodiment, the pressure oxidation may becarried out at a temperature of from about 115° C. to about 175° C. Thepressure oxidation may further be carried out under an oxygen partialpressure of from about 50 psig (345 kPa) to about 250 psig (1725 kPa).

Also according to the invention, there is provided a process for theextraction of a non-cuprous metal from a metal ore or concentrate,comprising the step of subjecting the ore or concentrate to pressureoxidation in the presence of oxygen and an acidic solution containinghalogen ions and a source of bisulphate or sulphate ions to form asolution of said non-cuprous metal, wherein said source of bisulphate orsulphate ions is selected from the group consisting of sulphuric acidand a metal sulphate which hydrolizes in said acidic solution. In thisspecification, M represents the metal being extracted, such as copper,zinc, nickel or cobalt.

The halogen concentration in the pressure oxidation filtrate, which isrecycled to the pressure oxidation step, is preferably maintained in therange of from about 8 g/L to about 20 g/L, preferably about 11 g/L toabout 14 g/L, and more preferably at about 12 g/L.

Reference is made to the use of chloride in the specification. However,it will be appreciated that the chloride could be substituted withbromide, if desired.

The term “non-cuprous metal” as used herein refers to a metal other thancopper.

The second leaching is preferably effected at a pH in the range of fromabout 1.3 to about 2.2. It has been found that this maximizes thesolution of base metal and minimizes the solution of iron. Morepreferably, the second leaching is effected in a pH range of from about1.6 to about 1.9.

The second leaching may be carried out at a temperature of from about20° C. to about 70° C., preferably, from about 35° C. to about 45° C.

For the second leaching, residence times of up to one hour or less, suchas 15 to 20 minutes, have been found to be adequate.

The raffinate may be split into a first portion comprising abouttwo-thirds of the raffinate and a second portion comprising aboutone-third of the raffinate and the first portion may be recycled to thesecond leaching and the second portion may be subjected to a secondarysolvent extraction to produce a secondary lixiviant and a secondaryraffinate. The secondary lixiviant may be used as extractant in thesolvent extraction of the leach liquor.

In another embodiment of the invention, the pressure oxidation iscarried out at a predetermined molar ratio of H³⁰/M, where H⁺ representsthe hydrogen ions in the acidic chloride solution and M represents themetal in the ore or concentrate, so that the pressure oxidation filtratecontains a first portion of the metal in the ore or concentrate and thebasic metal salt contains a second portion of the metal in the ore orconcentrate and further comprising the steps of separating the pressureoxidation filtrate and the basic metal salt; leaching the basic metalsalt in a second leaching step with an acidic sulphate solution todissolve the metal salt to produce a second metal solution and a solidresidue; and subjecting the pressure oxidation filtrate and the secondmetal solution to solvent extraction to produce concentrated metalsolution for electrowinning of metal therefrom.

The invention also extends to copper, zinc, nickel and cobalt wheneverproduced by the process according to the invention.

Further objects and advantages of the invention will become apparentfrom the description of preferred embodiments of the invention below.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a flow diagram of a hydrometallurgical copper extractionprocess according to one embodiment of the invention, which is suitablefor the treatment of high grade copper ores or concentrates.

FIG. 2 is a flow diagram of a hydrometallurgical copper extractionprocess according to another embodiment of the invention, which issuitable for the treatment of medium and lower grade copper ores orconcentrates.

FIG. 3 is a flow diagram of a hydrometallurgical copper extractionprocess according to a further embodiment of the invention, whichprovides for the extraction of zinc in addition to copper.

FIG. 4 is a flow diagram of a hydrometallurgical copper extractionprocess according to another embodiment of the invention, which providesfor the extraction of nickel in addition to copper.

FIG. 5 is a flow diagram of a hydrometallurgical process for theextraction of metals for a copper-nickel sulphide concentrate accordingto another embodiment of the invention.

FIG. 6 is a flow diagram of a hydrometallurgical process for theextraction of metals from a nickel-copper sulphide concentrate accordingto another embodiment of the invention.

FIG. 7 is a flow diagram of a hydrometallurgical process for theextraction of metals from a nickel laterite ore according to anotherembodiment of the invention.

FIG. 8 is a flow diagram of a hydrometallurgical process for theextraction of metals from a copper-zinc sulphide concentrate accordingto another embodiment of the invention.

DETAILED DESCRIPTION OF PREFERRED EMBODIMENTS

The process according to the invention is flexible enough to treat arange of copper concentrates in which the grade of copper varies fromlow, i,e. about 15% copper or less, to high grade, i.e. about 35% copperor more.

Broadly, the process comprises a pressure oxidation stage, anatmospheric leach stage, one or more solvent extraction stages and anelectrowinning stage. Different grades of concentrate require differenttreatment in the pressure oxidation stage, requiring different modes ofoperation. These modes of operation are termed Mode A and Mode B,respectively. In Mode A, which is effective when high grade copper oresare leached, copper is not leached in the pressure oxidation stage. InMode B, which is effective when medium and low grade copper ores areleached, copper is leached in the pressure oxidation stage.

Each of the two modes of operation will now be described in turn.

Process Mode A

FIG. 1 is a flow diagram of Mode A. The process comprises a pressureoxidation stage 12 in a pressure oxidation vessel or autoclave, anatmospheric leach stage 14, primary and secondary solvent extractantstages 16 and 18, respectively, and an electrowinning stage 20.

In the pressure oxidation stage 12, all copper minerals are converted tobasic copper sulphate, CuSO₄.2Cu(OH)₂. The treatment is carried out withoxygen in the presence of an acidic chloride solution. Oxygen, as wellas HCl and H₂SO₄ are introduced into the autoclave for this purpose. Thetemperature in the autoclave is about 130°-150° C. and the pressureabout 100-200 psig (1380 kPa). This is total pressure comprising oxygenpressure plus steam pressure. The retention time is about 0.5-2.5 hoursand the process is normally carried out in a continuous fashion in theautoclave. However, the process can also be carried out in a batch-wisefashion, if desired.

The solids content in the autoclave is maintained at about 12-25%, i.e.150-300 g/L solids as determined by the heat balance and viscositylimitations.

The slurry produced in the autoclave is discharged through a series ofone or more flash tanks 22 to reduce the pressure to atmosphericpressure and the temperature to 90°-100° C. The liquid part of theslurry is referred to as the product solution from the pressureoxidation stage 12 and is indicated by reference numeral 21.

The slurry from the flash tank(s) 22 is filtered, as shown at 24, andthe resultant filter cake is washed thoroughly to remove entrainedliquor as much as possible.

The pressure oxidation filtrate from the filtration 24 is recycled tothe pressure oxidation stage 12 but there is a small bleed of about 5%,as shown at 26. This bleed 26 is determined by the concentration of thesoluble metals in the ore or concentrate which may dissolve during thepressure oxidation stage 12. The bleed 26 is treated at 28 with lime toremove metals such as zinc and magnesium as solid residues, which arepresent in the copper concentrate, and to counteract buildup of thesemetals in the pressure oxidation circuit. The pressure oxidation circuitis the circuit from the pressure oxidation stage 12 to the flash tank(s)22 to the filtration 24 to the bleed 26 and back to the pressureoxidation stage 12. It is indicated by reference numeral 23.

The bleed 26 is subject to a solvent extraction, as shown at 27, priorto the bleed treatment 28. The solvent extraction 27 is carried out bymeans of a suitable organic extractant to remove copper from the bleed26. This solvent extraction is associated with the solvent extractionstages 16 and 18 and will be referred to again when the latter twosolvent extraction stages are described.

Prior to the pressure oxidation stage 12, the copper concentrate isfirst subjected to a regrind, as shown at 30, to reduce the particlesize to about 97% minus 325 mesh, which corresponds to P80 (80% passing)15 micron. The regrind 30 is carried out in solution recycled from thebleed treatment 28. Thus, the slurry from the bleed treatment 28 issubjected to a liquid/solid separation, as shown at 32, and the solutionis recycled to the regrind 30 and the zinc/magnesium bleed residue isdiscarded, as shown at 17.

The solution which is recycled to the regrind 30 is an alkaline chlorideliquor at about pH 10. Use of this liquor minimizes water input into thepressure oxidation circuit 23 which is important in maintaining heatbalance and in preserving the chloride solution in the pressureoxidation circuit 23 as much as possible.

As stated above, copper is not leached in the pressure oxidation stage12 but is converted to an insoluble basic copper salt. The feed solutionto the pressure oxidation stage 12, which is the leach liquor beingrecycled from the filtration 24 is indicated by reference numeral 25.Although there is copper present in the feed solution 25, there is noadditional copper leached, i.e. the process is operated so that thecopper concentration in the feed solution 25 to the pressure oxidationstage 12 is equal to the copper concentration in the product solution 21from the pressure oxidation stage 12. This is indicated as: Δ[Cu²⁺]=0.

The feed solution 25 to the pressure oxidation stage 12 contains about15 g/L Cu and 12 g/L Cl, together with about 30-55 g/L sulphuric acid.The acid is added in the form of make up H₂SO₄ (usually 93%). Theproduct solution 21 from the pressure oxidation stage 12 also containsabout 15 g/L Cu and 11-12 g/L Cl but is at about pH 3. There issubstantially no acid left in the product solution 21 as it is allconsumed in the pressure oxidation stage 12 to form the basic coppersalt.

As referred to above, the liquid pressure feed 25 to the pressureoxidation stage 12 is made up partly of recycled filtrate to which H₂SO₄is added. The immediate effect of adding the acid to the filtrate is toincrease the acidity of the filtrate which is fed to the autoclave forthe pressure leaching stage 12, but the most important effect,surprisingly, has been found to be that the addition of the acid, ormore specifically the sulphate ions, actually suppresses the oxidationof sulphur emanating from the concentrate in pressure oxidation stage12.

Typically the oxidation of sulphur that is experienced if no acid isadded is about 25-30% of the feed sulphur in the concentrate, as is thecase with the process described in U.S. Pat. No. 4,039,406. However, ifacid is added, it has been found that the sulphur oxidation to sulphateis reduced to about 5-10%. This improvement has substantial beneficialeffects on the hydrometallurgical extraction process. The oxidation ofsulphur to sulphate creates addition costs in several ways, such asadditional oxygen required for the reaction, additional reagent requiredto neutralize the acid so formed by the oxidation and provision must bemade for heat removal due to the oxidation of sulphur to sulphate whichis very exothermic. This actually limits the throughput of the autoclavein which the pressure leaching stage 12 takes place.

The chemistry of the reaction in the pressure oxidation stage 12 isbelieved to be altered by the addition of the acid as follows:

No acid addition:

3CuFeS₂+21/4O₂+2H₂O→[CuSO₄.2Cu(OH)₂]+3/2 Fe₂O₃+5S°  (1)

With acid addition:

3CuFeS₂+15/4O₂+H₂O+H₂SO₄→CuSO₄.2Cu(OH)₂+3/2 Fe₂O₃+6S°  (2)

In both reactions, the copper is precipitated in the form of a basiccopper salt, which has been found to comprise mostly basic coppersulphate.

In the first reaction it appears that the sulphate of the basic coppersulphate is supplied by oxidation of the feed sulphur in theconcentrate, whereas in the second reaction it appears to be supplied bythe sulphate ions in the acid which is added to the autoclave, thusobviating the need for the oxidation of sulphur to sulphate. Thus, inthe second reaction, there is a nett consumption of sulphate ions toform the basic copper salt. The amount of sulphuric acid needed tosuppress sulphur oxidation has been found experimentally to be about 25to 75 grams per liter, depending on the type of concentrate and thepercentage solids in the concentrate.

In actual test work, there is more sulphur oxidation than is predictedby either reaction. The first reaction predicts one sixth or 16.7% ofthe sulphur to be oxidized, whereas experimentally about 25%-30% isfound. With acid addition, experiments indicate that about 2-16% sulphuris oxidized to sulphate, rather than the zero oxidation that would bepredicted if the second reaction as written was the only reaction takingplace. Therefore, these reaction equations do not reflect exactly whatis happening in the pressure leaching stage 12 but are only anapproximation.

Chloride is conserved as much as possible in the pressure oxidationcircuit 23 but typically about 3-10% chloride is lost per pass into thesolid product at the filtration 24. Thus, the chloride must be made upby the addition of HCl or another source of chloride to provide 12 g/Lchloride in the feed solution 25. The chloride losses are minimized bythorough washing of the solids from the pressure oxidation stage 12 onthe filter 24. The amount of wash water is constrained by therequirement to maintain a water balance in the pressure oxidationcircuit 23. The only water loss from the circuit 23 is in the steam 29from the flashing step 22 and in the filter cake after the filtration24. Hence, the need to use the recycled solution from the bleedtreatment 28 to slurry up the concentrate in the grinding step 30, andthus minimize fresh water input from the concentrate to the pressureoxidation step 12.

It has been found to be advantageous to maintain at least 15 g/L Cu inthe product solution 21 from the pressure oxidation stage 12 so as tocounteract chloride loss in the form of solid basic copper chloride,CuCl₂.3Cu(OH)₂, which can occur if insufficient copper is present insolution to allow basic copper sulphate to form:

4CuCl₂+6H₂ 0→CuCl₂.3Cu(OH)₂+6HCl  (3)

This reaction can be counteracted by the addition of sufficient acidinto the autoclave during the pressure oxidation stage 12 to maintain atleast enough copper in solution to satisfy the stoichiometricrequirements for Cl as CuCl₂. For 12 g/L Cl in solution, thestoichiometric amount of Cu is:${\frac{63.5}{71} \times 12} = {10.7\quad {g/L}\quad {Cu}}$

Thus, 15 g/L Cu is a safe minimum to prevent a significant chloride lossin the form of the basic copper salt.

On the other hand, the copper concentration in the product solution 21from the pressure oxidation stage 12 should be kept as low as possibleto counteract the formation of CuS by the reaction of elemental sulphurwith aqueous copper sulphate. This reaction can occur during thepressure oxidation stage 12 or in the slurry after discharge from theautoclave but before the filtration step 24:

3CuSO₄(aq)+4S°+4H₂O→3CuS(s)+4H₂SO₄  (4)

This reaction is particularly undesirable because CuS is insoluble inthe dilute acid conditions of the atmospheric leaching stage 14. Thus,the copper is not recovered and this results in the loss of copper tothe final residue.

To counteract the formation of CuS it is necessary to keep the copperconcentration in the product solution 21 as low as possible, i.e. below30 g/L for some concentrates. The tendency to CuS formation isapparently related to the type of concentrate being treated, with themedium to high grade concentrates being more susceptible to CuSformation. Thus, although a high copper concentration in the productsolution 21 does not present a problem with the low grade concentrates,it cannot be tolerated with the higher grade concentrates.

As is known to date, high grade concentrates, i.e. above 35% copper, arebest treated to produce as low a copper concentration in the productsolution 21 as possible, i.e. below 25 g/L Cu.

Given the need to maintain at least 15 g/L Cu in solution in thepressure oxidation circuit 23, there is an optimum range of copperconcentration of from 15 to 25 g/L Cu for high grade concentrates. Withmedium grade concentrates, the upper limit can be stretched considerablyand for low grade ore, the copper concentration does not play asignificant role.

The copper concentration in the pressure oxidation filtrate 29 can becontrolled simply by adding the required amount of acid into the feedsolution 25 to the pressure oxidation stage 12. More acid results in ahigher copper concentration due to the dissolution of the basic coppersulphate:

CuSO₄.2Cu(OH)₂(s)+2H₂SO₄→3CuSO₄(aq)+4H₂O  (5)

The addition of about 1 g/L acid results in an increase in copperconcentration of about 1 g/L. The actual concentration of acid requiredis determined empirically by comparing the assays of feed solution 25 tothe pressure oxidation stage 12 and the product solution 21 from thepressure oxidation stage 12 to satisfy Δ[Cu²⁺]=0. The volume of solutionin the circuit 23, however, is determined by the heat balance.

The percentage by weight of solids in the feed of copper concentrateslurry to the pressure oxidation stage 12 can be varied at will. Theweight of concentrate solid fed to the pressure oxidation stage 12 isdetermined by the amount of copper to be recovered. The weight of thesolution is determined mainly by the heat balance in the pressureoxidation stage 12.

The desired operating temperature in the pressure oxidation stage 12 isabout 150° C. and the heat must be supplied largely by the heat ofreaction of the sulphide minerals with the high pressure oxygen in theautoclave. For high grade concentrates, such as will be treated by theProcess Mode A currently being described, this means a relatively lowS/Cu ratio and thus a smaller heat production per tonne of coppertreated in the autoclave. Much of the heat evolved is due to oxidation,not of copper, but of the other two main elements in the concentrate,iron and sulphur. If the grade of the concentrate is high, then theratio of S/Cu and Fe/Cu is low, hence a lower heat production.

To reach operating temperature from a starting temperature of say 50° to80° C., which is typical for the pressure oxidation filtrate 29 which isrecycled after the filtration 24, it is necessary to control the amountof water that must be heated, since this is the main heat sink in thepressure oxidation stage 12. It is impractical to cool or heat theslurry inside the autoclave by indirect means, such as by means ofheating or cooling coils, because of rapid scale formation on allsurfaces, particularly heat exchangers, leading to very poor heattransfer characteristics. Direct heating or cooling by injection ofsteam or water is also impractical due to water balance considerations.Therefore, it is required that the heat balance be maintained bybalancing heat production from reaction heat with the heat capacity ofthe feed materials, i.e. the feed solution 25 being recycled and theconcentrate slurry. The main variable that can be controlled here is thevolume of the feed solution 25. This is one of the distinguishingfeatures between Modes A and B. In Process Mode B, still to bedescribed, the heat evolution is much greater, expressed as heat pertonne of copper product. Therefore, it is possible to use more solutionvolume in the feed 25 to the pressure oxidation stage 12.

Once the solution volume is fixed, the acidity of the solution can bedetermined, since the total mass of acid is determined by the need tomaintain Δ[Cu²⁺]=0. Typically, for a high grade concentrate, about 35-55g/L acid will be required.

It has been found to be beneficial to add small concentrations ofcertain surfactants which change the physical and chemicalcharacteristics of liquid elemental sulphur (S°) in the autoclave duringthe pressure oxidation stage 12. Surfactants such as lignin sulphonateand quebracho added to the pressure oxidation feed solution 25 in smallamounts, i.e. 0.1 to 3 g/L can reduce the viscosity of the liquidsulphur and also change the chemistry in the autoclave.

Additions of surfactants can reduce the sulphur oxidation in ways thatare not well understood, but are beneficial to the process. It isbelieved that this is due to lower viscosity, resulting in loweredtendency for liquid sulphur and solids to be held up within theautoclave, thus reducing the retention time for these materials, andhence the reduced tendency for sulphur oxidation to occur.

Also it has been found that more complete reaction of the copperminerals takes place if surfactants are added, apparently because oflower viscosity sulphur, which does not “wet” unreacted sulphideminerals, and thus allows the desired reaction to proceed to completion.

Reaction (5) describes how adding sulphuric acid to the pressureoxidation feed 25 will control the copper concentration in the pressureoxidation filtrate 29. The overall reaction for the pressure oxidationwith sulphuric acid addition for a chalcopyrite ore is given by reaction(2) above.

A similar reaction can be written using CuSO₄ as the source of sulphideions instead of H₂SO₄:

3CuFeS₂+15/4O₂+3H₂O+3/2 CUSO₄→3/2 CuSO₄.2Cu(OH)₂+3/2 Fe₂O₃+6S°  (6)

It is noteworthy that there are 3/2 moles of sulphate required as coppersulphate in reaction (6) compared to one mole of sulphuric acid inreaction (2). Therefore, if CuSO₄ is to be used as the source ofsulphate ions instead of sulphuric acid, it is necessary to use 1.5times as many moles of CuSO₄. To take this into account, the inventorhas developed the concept of Excess Sulphate Equivalent, which allowsthe calculation of how much acid to add to the pressure oxidation feedsolution 25 in order to achieve a target copper concentration and stilltake into account reaction (6).

By taking reaction (6) into account, it is possible to calculate “apriori” the amount of acid required for constant copper concentration inthe pressure oxidation filtrate 29. The concept of Excess SulphateEquivalent is helpful:

Excess Sulphate Equivalent is equal to the sulphate available in thepressure oxidation feed solution 25 for formation of basic coppersulphate during the pressure oxidation stage 12. The sulphate availableis that which is in excess of a defined Base Level of CuSO₄ and CuCl₂.

Base Level of CuSO4 and CuCl₂ is sufficient to support chloride insolution at 12 g/L in the form of CuCl₂ and, in addition, about 4.3 g/LCu as CuSO₄. The concentration of CuCl₂ corresponding to 12 g/L chloridein solution is 134.5/71*12=22.7 g/L CuCl₂, which contains 10.7 g/L Cu insolution. The additional 4.3 g/L copper therefore means a total of 15g/L copper combined as CuCl₂ and CuSO₄ in the Base Level.

Sulphate available is then the total sulphate as CuSO₄ less the BaseLevel. For instance, if the total copper concentration is 28 g/L in thepressure oxidation filtrate 29, then the sulphate available is 28−15=13g/L Cu * 98/63.5=20 g/L H₂SO₄ as available sulphate from CuSO₄.

Excess Sulphate Equivalent (ESE) is then calculated from the availablesulphate from CuSO₄ by dividing by 1.5:

ESE={Available Sulphate as CuSO₄}/1.5

Thus, in the example of 28 g/L total copper concentration or 20 g/Lavailable sulphate from CuSO₄, there is 20/1.5=13.3 g/L ESE from CuSO₄.

Finally, if the target free acid equivalent is, say, 52 g/L H₂SO₄ in thepressure oxidation feed solution 25, then the amount of acid required is52 less the ESE (13.3 g/L) or 38.7 g/L H₂SO₄. This is the amount thatmust be added to the feed solution 25 to the pressure oxidation stage 12to produce a constant copper concentration in the pressure oxidationfiltrate 29, i.e. the Base Level of 15 g/L Cu.

Other reactions can be written using Fe₂(SO₄)₃ and ZnSO₄ as the sourceof sulphate ions instead of H₂SO₄. In the case of ZnSO₄, the zinc isassumed to hydrolyze to basic zinc sulphate, ZnSO₄.3Zn(OH)₂, which is abasic salt of Zn analogous to basic copper sulphate. These reactions aregiven below as reactions (7) and (8).

3CuFeS₂+15/4O₂+2H₂O+1/3 Fe₂(SO₄)₃→CuSO₄.2Cu(OH)₂+11/6Fe₂O₃+6S°  (7)

3CuFeS₂+15/4O₂+13/3H₂O+4/3 ZnSO₄→CuSO₄.2Cu(OH)₂+6S°+Fe₂O₃+1/3{ZnSO₄.3Zn(OH)₂.4H₂O}  (8)

The solids from the pressure oxidation stage 12 after the filtration 24,are treated in the atmospheric leaching stage 14 at about pH 1.5 to pH2.0 using raffinate from the primary leaching stage 16, which is acidic,to dissolve the basic copper sulphate. The leaching 14 takes place at atemperature of about 40° C. for a retention time of about 15-60 minutes.The percentage solids is typically about 5-15% or about 50-170 g/L,although it is possible to operate the process outside this range.

During the atmospheric leaching stage 14, the basic copper saltsdissolve almost completely with very little of the iron present in theconcentrate going into solution.

Typically, the leach liquor 33 produced after the liquid/solidseparation 34 contains about 10-20 grams per litre copper, depending onthe percentage solids feed to the leach 14, with 0.1-1.0 g/L iron andabout 0.1-1.0 g/L chloride. Much of this iron and chloride are derivedfrom the feed raffinate 37 rather than the solids from pressureoxidation, i.e. they are recycled. Typically about 0.1-0.2 g/L iron andchloride dissolve per pass.

The copper extraction has been found to be about 95-98% based on theoriginal feed to the pressure leaching stage 12. Iron extraction tosolution has been found to be less than about 1%.

The slurry 31 from the atmospheric leaching stage 14 is difficult if notimpossible to filter, but settles well. In view of the need to wash theleach solids very thoroughly, the slurry 31 is therefore pumped to acounter current decantation (CCD) wash circuit, symbolically indicatedas a solid/liquid separation 34 in FIG. 1. In the CCD circuit 34, thesolids are fed through a series of thickeners with wash water added inthe opposite direction. By this method, the solids are washed andentrained solution removed. About 3 to 5 thickeners (not shown) arerequired with a wash ratio (water to solids) of about 5 to 7 to reduceentrained liquor down to less than 100 ppm Cu in the final residue.

The thickener underflow from the last thickener is the final residuestream 35 at about 50% solids. This can be treated for the recovery ofprecious metals, such as gold and silver, or sent to tailings. Preciousmetals may be recovered by known methods, such as cyanidation. The mainconstituents of the stream 35 are hematite and elemental sulphur, whichmay be recovered by flotation if market conditions warrant.

The thickener overflow from the first thickener is the product solution33 which is fed to the primary solvent extraction stage 16, as shown.This solution contains about 12 g/L Cu, 1 g/L Cl and 0.5 g/L Fe.

The optimum copper concentration is determined by the ability of thesolvent extraction stage 16 to extract the maximum copper from thesolution 33. Since a fraction of about one-third of the raffinate fromthe solvent extraction stage 16 is eventually neutralized, it isimportant to minimize the copper content of this raffinate.

Solvent extraction performs best on dilute copper solutions due to thefact that a concentrated copper solution results in a higher acidconcentration in the raffinate which tends to lower extractionefficiency. More concentrated solutions are, however, cheaper to treatfrom a capital cost point of view, since the volume is less. Above acertain point, though, the increased concentration does not reduce thesize of the solvent extraction unit, since (i) there is a maximumorganic loading and (ii) aqueous volume is generally kept equal toorganic volume for mixing purposes by means of aqueous recycle.Therefore, the total volume of organic extractant and aqueous solutionis only determined by the volume of organic extractant. The maximumorganic loading and hence volume of organic is determined by theconcentration and characteristics of the particular organic solventselected. For the typical solvent, e.g. LIX™ reagent from HenkelCorporation, the maximum loading per pass at 40% volume concentration indiluent is about 12 g/L Cu. Therefore, the product solution 33 alsoshould contain about 12 g/L Cu.

The copper is extracted from the product solution 33 from the CCDthickener overflow in two stages of extraction in the primary solventextraction stage 16 to produce a raffinate 37 with about 20 g/L freeacid and about 0.3 to 1 g/L Cu. Most of this raffinate 37 is recycled tothe atmospheric leaching stage 14 but about 25 to 30% is surplus to theacid requirements of the atmospheric leaching stage 14 and must beneutralized. This surplus 121 is split off as shown at 36 andneutralized.

The neutralization is effected in two stages to maximize copper recoveryand to prevent possible environmental problems with the neutralizationresidue due to copper content, i.e. the unrecovered copper from theraffinate 37 will precipitate upon neutralization and can thenre-dissolve later, in a tailing pond, for example.

The first stage neutralization takes place at pH 2 to pH 3, as shown at38, using limerock, which is very economical as a reagent, compared withlime. The neutralization product is filtered at 40 and the resultantsolids are washed with water from the external source 45. The solids,which are mainly gypsum and iron hydroxides, are discarded, as shown at41.

The filtrate 39 is sent to the secondary solvent extraction stage 18 forthe recovery of residual copper values. The secondary solvent extraction18 benefits from the primary neutralization 38 and results in a very lowcopper concentration in the secondary raffinate 43, typically from about0.03 to 0.06 g/L Cu.

As indicated by the broken lines in FIG. 1, the secondary solventextraction stage 18 uses the same organic extractant as the primarysolvent extraction circuit 16. This is also tied in with the solventextraction 27 of the pressure oxidation filtrate bleed 26. The organicextractant which is washed at 42 with wash water 122 from an externalsource 45, and stripped at 44 is recycled to the secondary solventextraction stage 18 and then passes to the primary extraction stage 16.The stripped organic 125 is split to pass a portion thereof to thesolvent extraction 27. The raffinate from the solvent extraction 27 isadded to the loaded organic 123 from the solvent extraction 16 prior tothe wash 42. The wash water 47 from the wash 42 is passed to thepressure oxidation filter 24, to serve as a feed wash water onto thefilter 24. The resultant wash filtrate is added to the pressureoxidation filtrate 29, thus recovering the copper and chloride contentfrom the solvent extraction wash water (47).

The raffinate 43 from the secondary solvent extraction stage 18 isneutralized again in a secondary neutralization stage 46, this time atpH 10 and filtered at 48 to remove all dissolved heavy metals, producinga solution 51 which is used as wash water in the CCD circuit 34 forwashing the final leach residue 35. The solid residue from thefiltration 48 is discarded, as shown at 53.

Stripping of the loaded and washed organic at 44 is effected by means ofspent acid or electrolyte 55 from the electrowinning stage 20 to obtaina pure copper sulphate solution or pregnant electrolyte 57 which is thenpassed to the electrowinning stage 20 for electrowinning in the usualway.

It can be seen that all solution streams in the process are thusrecycled and there are no solution effluents from the process. Onlysolid residues are discarded from the process.

Process Mode B

FIG. 2 is a flow diagram of Mode B. The same reference numerals are usedto indicate stages or steps in the process which correspond with thosein the previous embodiment of FIG. 1. For example, the pressureoxidation stage is again indicated by 12, the atmospheric leach stage by14, the electrowinning stage by 20, the flash tank(s) by 22, thepressure oxidation filtration by 24, the bleed treatment of the pressureoxidation filtrate 29 by reference numeral 28, the grinding stage byreference numerals 30 and the CCD wash circuit by reference numeral 34.

In this mode of the process, the pressure oxidation 12 is carried outboth to oxidize and to leach into solution most of the copper containedin the feed concentrate. Typically about 85-90% of the copper is leachedinto the solution, with only about 10-15% being left in the residue asthe basic copper sulphate.

The conditions of the pressure oxidation stage 12 in the autoclave aresimilar to those in Process Mode A except that the percentage solids islower, i.e. 150-225 g/L.

In this mode of the process, Δ[Cu²⁺] is typically 30 to 40 g/L Cu, i.e.the copper concentration is greater in the product solution 21 from thepressure oxidation stage 12. The feed solution 25 to the pressureoxidation stage 12 typically contains 10-15 g/L Cu and 12 g/L Cl,together with about 20 to 30 g/L sulphuric acid.

In this mode, no sulphuric acid is added to the pressure oxidation stage12 from an external source, as is the case with the FIG. 1 embodiment.In this mode, the acid is obtained from recycle in the process, i.e. bythe recycle of the pressure oxidation filtrate 29. The product solution21 from the pressure oxidation stage 12 contains about 40 to 50 g/L Cuand 11 to 12 g/L Cl at about pH 2 to 2.5.

The copper leached into the product liquor 21 from pressure oxidationstage 12 must be controlled so as to obtain the desired distribution ofcopper between liquor (85 to 90%) and residue (10 to 15%). Thisdistribution results in a small but important amount of basic coppersulphate solids in the leach residue. The pH is a convenient indicatorof the presence of basic copper sulphate, since it is a buffering agent.With strong copper sulphate concentration in solution, a pH range of 2to 2.5 indicates basic copper sulphate. Below pH 2 almost all the basiccopper sulphate will be dissolved, whereas above pH 2.5, too much basiccopper sulphate is formed and insufficient copper is likely to be foundin the solution 21.

The primary method of control is the amount of acid in the feed liquor25 to the pressure oxidation stage 12. The acid level in turn iscontrolled by the degree of neutralization of the raffinate from solventextraction of the pressure oxidation filtrate 29 raffinate describedbelow. Usually, about 25 to 50% of the acid must be neutralized,depending on the amount of acid that is required.

The acid generated during the pressure oxidation stage 12 varies fromone concentrate to another and according to conditions employed. If theconcentrate produces a large amount of acid during the pressureoxidation stage 12, then the feed solution 25 will need less acid toachieve the desired result. The minimum copper (from concentrate feed)that should go to liquor 21 is about 10%. Below 10%, the pH dropssufficiently low so that iron concentrations in the pressure oxidationfiltrate 29 increase rapidly. Normally, iron is about 10 to 50 ppm, butif pH is below 2 and basic copper sulphate in residue disappears, theniron can increase to above 1 g/L fairly quickly. This is undesirablebecause there are several impurity elements such as As and Sb which areonly removed from solution simultaneously with iron hydrolysis.Therefore, absence of iron in solution is a good guarantee of lowimpurity content in the pressure oxidation filtrate 29. Iron is also animpurity itself that must be avoided in the electrowinning circuit 20 asfar as possible.

There is another factor, however, which places a maximum on Cu insolution. It has been found surprisingly that certain concentratesactually leach more completely if the copper concentration is lower.This is believed to be due to either formation of secondary CuS, asdescribed above, or to some other phenomenon related to poor oxidationcharacteristics of the primary mineral, chalcopyrite, in high copperconcentration solutions. It is found that elemental 10 sulphur, producedduring the reaction in the pressure oxidation stage 12, can coat oractually encapsulate unreacted chalcopyrite particles and hinder theaccess of reagents. This results in poor copper recovery. The phenomenonis apparently accentuated by high Cu levels in solution. It can beovercome or mitigated by the use of surfactants, as described above. Theproblem is more severe with some concentrates, particularly high grade,than others. Therefore, for these concentrates it is desirable to limitthe copper concentration in the pressure oxidation filtrate (i.e.greater than about 95%) over all. To do this, it is necessary to have asubstantial proportion of the copper as basic copper sulphate, i.e. insolid residue from the pressure oxidation stage 12 rather than thepressure oxidation filtrate. Typically, 20-40% of copper may report tosolids, if necessary, to keep the copper concentration low enough toobtain high copper recovery.

Higher grade concentrates exhibit the problem of low copper recoverywith high copper in solution. Therefore, an increasing proportion ofcopper must report to solids as the grade increases. Tests with threedifferent concentrates illustrate this relationship:

H*/Cu Cu Distribution % Total Conc. # % Cu Molar PO liquor PO residuerecovered 1 41 0.55 0 100 97.3 2 28 0.70 63 37 95.7 3 22 0.96 85 15 94.7

The H⁺/Cu molar ratio refers to H⁺ in the feed acid and Cu in the feedconcentrate. The H⁺ in the feed acid is taken to be all the protonsavailable on complete dissociation of the acid even if under existingconditions the acid is not completely dissociated. The H⁺ shown in thetable is optimum level found by experiment to give the best results.

For concentrate #1, which was a high grade concentrate, the processchosen is Mode A, where all of the copper reports to the leach liquorand Δ[Cu²⁺]=0. The H⁺/Cu ratio is that found which was necessary byexperimentation to give the desired result of Δ[Cu²⁺]=0.

For concentrate #2, a medium grade concentrate, Mode B was chosen, butwith a substantial amount of the copper reporting to the solid basiccopper sulphate. This was achieved by keeping the H⁺/Cu ratio low enoughso that not all of the copper dissolved into the liquor.

For concentrate #3, a low grade concentrate, Mode B was also chosen butin this case, the minimum amount of copper reported to the residue, byadjusting the H⁺/Cu ratio high enough.

The residue from the pressure oxidation stage 12 is leached 14 withraffinate 37 returning from the solvent extraction 16 which is diluteacid, at 3-10 g/L H₂SO₄. Since most of the copper from the pressureoxidation stage 12 reports to the pressure oxidation filtrate 29 andonly a small fraction of the pressure oxidation residue, the resultantleach liquor 31 from the atmospheric leach 14 is quite dilute in copper.In turn, this produces a dilute raffinate 37 from the solvent extraction16. Typically, the atmospheric leach liquor 31 is 3-7 g/L Cu and 0.2 to0.5 g/L Fe.

The slurry resulting from the atmospheric leaching stage 14 is difficultto filter, as was the case with Mode A. Good liquid/solid separation andwashing, however, can be achieved as before using a series of thickenersin a CCD arrangement 34. Wash water 51 is provided by raffinate from thesolvent extraction 16, which is neutralized, as indicated at 46. This issimilar as in Mode A. The only major difference is the lower tenor ofthe solution 33 and the reduced volume.

The solution 33 produced by the atmospheric leaching stage 14 issubjected to the solvent extraction 16. The copper containing solution29 from the pressure oxidation stage 12, is subject to a solventextraction stage 50. There are, therefore, two solvent extractionoperations, i.e. 16 and 50, treating two different streams of liquor 33and 29, respectively. It is a feature of the process according to theinvention that the organic extractant used for effecting the solventextraction operations is common to both solvent extractions 16 and 50.

As shown in FIG. 2, the stripped organic 125 coming from the commonstripping operation 44 is first introduced into the solvent extractioncircuit 16, which has the weakest copper concentration in the aqueousfeed stream 33 and therefore needs the organic extractant to be as lowas possible in loading to be effective.

The loaded organic 126 from solvent extraction 16 is then sent to thesolvent extraction 50 where it contacts the higher copper concentrationliquor 29. It is not necessary for the solvent extraction 50 to achievea high extraction ratio because the raffinate 63 from this extraction isrecycled to the pressure oxidation stage 12, a shown. On the other hand,the raffinate 37 from the solvent extraction 16 is only partly recycledand part is neutralized 46 to remove excess acid from the circuit.Therefore, it is more important to achieve high copper recovery from thesolvent extraction 16.

The raffinate 37 from the solvent extraction 16 is split at 36 as inMode A, with about one-third 121 to the neutralization 46 and two-thirds120 recycled to the atmospheric leach stage 14. An important differencefrom Mode A is that the raffinate 37 from solvent extraction 16 issufficiently low in copper, i.e. below 100 ppm, so that it is notnecessary to have a secondary solvent extraction stage beforeneutralization 46, as was the case in Mode A. This is due to the lowercopper concentration and solution volume, allowing the solventextraction 16 to be more efficient.

The loaded organic 65 produced by the two solvent extraction operations16, 50 in series, is washed in two stages in counter current fashionwith dilute acidic aqueous solution 122, as shown at 42. This isprimarily to remove entrained aqueous solution from the loaded organic65 and in particular to reduce the chloride content before the organicgoes to stripping at 44. The amount of wash water required is about 1-3%of the organic volume. The resultant wash liquor 47 produced is recycledto the pressure oxidation stage 12.

The washed organic 69 is stripped at 44 with spent electrolyte 55 fromthe electrowinning stage 20 to provide a pure copper solution orpregnant electrolyte 57 for electrowinning in the usual way.

The raffinate 63 is split at 70 in two portions 72, 74 as determined bythe required molar ratio of H⁺/Cu. The portion 72 is recycled to thepressure oxidation stage 12. The portion 74 is neutralized at pH 2 withlimerock at 76 and filtered 78. The solid residue is washed anddiscarded, as shown at 80. The filtrate 82 is recycled with the portion72 to form the feed solution 25 to the pressure oxidation stage 12.

A novel feature of the process, therefore, is the use of a commonorganic to extract copper from two separate aqueous feed liquors. Thisprovides considerable economies in lower capital and operating costs inthe solvent extraction circuits. Also, it allows for the use of copiousamounts of water in the atmospheric leaching CCD circuit, so that goodwashing can be achieved on the final residue and yet still recovercopper from such a dilute liquor.

It has been found that the degree of sulphur oxidation that occurs inthe pressure oxidation stage 12 is highly dependent on the type ofconcentrate, such as grade and mineralogy of the concentrate beingtreated, as well as the conditions of the pressure oxidation stage 12.Certain concentrates exhibit considerably higher sulphur oxidation, i.e.oxidation of the sulphur in the concentrate to sulphate, and the effectis particularly marked with the low grade concentrates with less thanabout 28% Cu by weight. The inventor has found that the significance ofthis variation is not so much the copper grade itself but thecopper/sulphur ratio in the concentrate. The main impurity elements in acopper concentrate are iron and sulphur due to the fact that copper oresare generally composed of chalcopyrite together with other minerals,particularly pyrite FeS₂ or pyrrholite FeS.

Process Mode B deals with the problem of excess sulphur oxidation in thepressure oxidation stage 12 when lower grade concentrates are used bydeliberately dissolving 90% of the copper and minimizing the formationof basic copper sulphate. The reaction for chalcopyrite is:

CuFeS₂+5/4O₂+H₂SO₄→CUSO₄+1/2Fe₂O₃+2S°+H₂O  (6)

The filtrate 29 from the pressure oxidation stage 12 thus contains highlevels of copper sulphate and copper chloride and this is treated in thesolvent extraction stage 50 to produce a pure copper sulphate solutionwhich goes to the electrowinning stage 20.

With reference to FIG. 3, a hydro-metallurgical process for extractionof zinc in addition to copper is shown. The same reference numerals areused to indicate stages or steps in the process which correspond withthose in the previous embodiments.

The concentrate is re-ground 30 as in the case of the previousembodiments.

The pressure oxidation of a mixed zinc-copper concentrate is carried outin similar fashion as for the concentrate containing only copper as inFIG. 2.

Zinc oxidizes as readily or more readily than copper does and is morelikely to report to the leach liquor 29 as opposed to the pressureoxidation residue. This is because zinc hydrolyzes less readily as basiczinc sulphate than copper does, i.e. at higher pH.

Recovery of copper or zinc is not hampered by high solution tenorsapparently as was found for high grade copper concentrations. Therefore,it is possible to have most of the copper and zinc report to thepressure oxidation filtrate 29, i.e. as in Process Mode B. Sulphuroxidation is low, so that the amount of acid generated within thepressure oxidation stage 12 is low. Hence, to obtain a high H⁺/Cu ratio,it is necessary to recycle virtually all of the acid from the solventextraction stage 12 with minimal neutralization. The feed acid may be ashigh as 75 g/L H₂SO₄ with about 10 g/L Cu, 5 g/L Zn and 12 g/L Cl.

The pressure oxidation filtrate 29 will contain both zinc and copper insubstantial concentrations dependent on the feed concentratecomposition. For a concentrate with 20% Cu and 5% Zn, the pressureoxidation filtrate 29 may contain about 50 g/L Cu, 15 g/L Zn and 12 g/LCl.

The pressure oxidation residue is leached 14 in the same way usingraffinate 37 from the solvent extraction 16 as shown, producing a mixedCu-Zn solution for feed to the solvent extraction circuits. Zinc isextracted first and then copper.

There are two aqueous streams to be treated by solvent extraction as inProcess Mode B for copper concentrates. The pressure oxidation filtrate29 contains high tenors of Cu and Zn, whereas the atmospheric leachsolution 33 is weak in both elements.

The novel arrangement outlined for the solvent extraction circuit as forthe embodiments described above, is continued for the zinc solventextraction, that is, the weak liquor is contacted first with organicextractant followed by the strong aqueous liquor. In this case, thereare two circuits, one for zinc and one for copper.

It is possible to extract copper first followed by zinc, depending onthe choice of organic extractant and its relative affinity for the twoelements. The applicant has found that satisfactory results can beobtained by using DEHPA (diethyl-hexyl phosphoric acid) as the firstextractant, which is selective towards zinc over copper. Therefore, twoDEHPA extractions 100 and 102 are done, the first extraction 100 is onthe weak liquor 33 and the second extraction 102 is on the strongerliquor 29 from the pressure oxidation stage 12, to recover zinc andleave the bulk of the copper in solution.

The zinc extraction by DEHPA is hampered by poor extractioncharacteristics in the presence of high acid concentrations. Inpractice, this means that the extraction effectively stops at about pH1.4 or about 7-10 g/L H₂SO₄. To deal with this problem, an interstageneutralization 104 at pH 2 is included for the zinc solvent extraction.Thus, the zinc solvent extraction occurs in two stages, i.e. the stage102 and a second stage 103 with the neutralization 104 in between. Eachstage 102, 103 will extract only 5-7 g/L zinc before being stopped bythe resultant acid concentration in the raffinate. By using interstageneutralization 104, the total zinc extraction can be increased to 10 g/LZn or more. The raffinate 97 from the first extraction stage 102 isneutralized to about pH 2 to 2.5 at 104 with inexpensive limerock(CaCO₃) to produce gypsum solids which are filtered off at 98 anddiscarded. The filtrate 99 is then fed to the second solvent extractionstage 103. The feed to the second stage is typically 10 g/L Zn and 50g/L Cu at a pH of 2 to 2.5. After extraction, the second raffinate 124is typically 5 g/L Zn, 50 g/L Cu and 8 g/L acid.

For the solvent extraction circuit 16, zinc concentrations are lowenough so that this does not present a problem.

The optimum zinc content of the pressure oxidation filtrate 29 isdetermined largely by the ability of the zinc solvent extraction circuitto extract the zinc. Due to the fact that zinc is extracted quite weaklyby the available extractants (e.g. DEHPA), there is a maximum of about5-7 g/L Zn that can be extracted before the reaction stops due to acidbuildup in the raffinate. Further extraction requires neutralization ofthe acid. With interstage neutralization it is possible to extract muchhigher levels of Zn, however, the interstage neutralization removessulphate from the circuit which must be replaced either by sulphuroxidation or adding fresh acid to the pressure oxidation circuit 23.

One inter-neutralization stage is likely to be compatible with sulphatebalance, therefore it is preferable to keep the Δ[Zn²⁺], which is thezinc concentration in the pressure oxidation filtrate 29 minus the zincconcentration in the recycled raffinate 72, to about 10 g/L. Thus, ifthe feed acid to pressure oxidation recycled as raffinate 72 fromsolvent extraction contains 5 g/L Zn, then the product filtrate 29 frompressure oxidation should contain about 15 g/L Zn. This restriction onΔ[Zn] distinguishes the process for Zn compared to Cu. The greaterextraction ability of Cu solvent extraction means that good extractionof Cu can be achieved with much higher acid levels, up to 75 g/L H₂SO₄in raffinate compared to only about 7-10 g/L for Zn. Hence Cu can beextracted from 50 g/L Cu feed streams.

After extraction, the loaded organic 106 from the Zn (DEHPA) circuitcontains some Cu, as a result of imperfect selectivity of the DEHPAextractant towards Zn, and simple entrainment of the strong Cu liquor.Typically the Zn/Cu ratio in the loaded organic 106 from Zn solventextraction is about 150 to 300:1. If not removed, all of the Cu will bestripped along with the Zn during solvent stripping 114, and thus willbe stripped into the Zn pregnant electrolyte 120 which is fed to Znelectrowinning 118. Zn electrowinning requires a very pure pregnantelectrolyte if it is to produce satisfactory (pure) Zn cathode atreasonable current efficiency. The Zn/Cu ratio must be about 100,000:1in pregnant electrolyte. Therefore it is essential to remove almost allof the Cu either from the loaded organic 106 or later from the pregnantelectrolyte before electrowinning. It is much easier to purify theloaded organic 106.

To remove this copper, several washing or treatment stages 106, e.g. 3to 10, typically 5, are needed. Washing is done with dilute acidifiedzinc sulphate aqueous solution. The wash stages are arranged in series,i.e. the treated organic exiting from the first wash stage enters thesecond wash stage and so through all the other stages until the organicexits the last stage. Some zinc is washed out with the copper,therefore, it is necessary to minimize the amount of wash water addedand make use of several wash stages arranged in counter current fashioninstead.

The resultant wash liquor 110 produced is recycled to the atmosphericleach circuit for recovery of copper and zinc values.

After washing, the organic stream 112 from the DEHPA extraction is readyfor stripping 114 with spent electrolyte 116 from a zinc electrowinningcircuit 118. This produces a pregnant electrolyte 120 for electrowinningzinc at high current efficiency.

After the stripping 114 the extraction solvent is further stripped 131for removal or iron prior to recycling of the extractant to the solventextraction 100. The stripping 131 is effected with HCl makeup solution133 which is introduced into the pressure oxidation stage.

The raffinates 122, 124 from the zinc extractions with DEHPA are eachextracted with a selective copper extractant, such as LIX™, in solventextractions 16 and 50, respectively.

The design of these two circuit 16, 50 is similar as in Process Mode Bwith a common organic used first in the solvent extraction 16 and thenin the solvent extraction 50. The loaded organic is then washed andstripped as before as shown at 42 and 44, respectively.

Neutralization requirements in the solvent extraction 50 circuit arefound to be low because of the prior neutralization in the zinc circuit.

The raffinates from the LIX™ extractions are recycled as before back tothe pressure oxidation stage 12 and the atmospheric leach stage 14,respectively.

With reference to FIG. 4, a hydrometallurgical extraction process forrecovery of nickel in addition to copper is shown.

The same reference numerals are used to indicate stages or steps in theprocess which correspond with those in the previous embodiments.

For nickel-copper concentrates, the process is very similar as for zinc,except that the available solvent extraction agents are all lessselective toward nickel than copper. Therefore, the nickel solventextraction circuits 130, 132 both are positioned after the respectivecopper solvent extraction circuits, 16, 50, respectively.

The loaded nickel extractant 135 from the solvent extraction 132 issubject to a wash 137 and then stripped 139 before being recycled to thesolvent extraction 130. The stripping 139 is effected with spentelectrolyte from the nickel electrowinning 140.

In addition, nickel extraction is sufficiently weak that in situneutralization with ammonia, for example, is required, as indicated at134 and 136, respectively. The ammonia must be recovered from therespective raffinates by a lime boil process 138, for example, andrecycled.

It has been found that there is a limit to the amount of sulphuroxidation that can be accommodated by the process Mode B. If the sulphuroxidation is high enough and sufficient acid is generated duringpressure oxidation, there will be a surplus of acid left over afterpressure oxidation, even if no acid is added to the feed, such as in theform of acidic raffinate. In this situation, not only will all thecopper in the concentrate be converted to dissolved copper sulphate, butalso some of the iron in the concentrate will be solubilized by thesurplus acid, e.g. as ferric sulphate.

It is desirable that iron in the concentrate report to the pressureoxidation residue as stable hematite, Fe₂O₃, and not to the solution,where it must be separated from the copper. Typical concentrates have anFe:Cu ratio of at least 1:1, and therefore the efficient and completeelimination of Fe at an early stage is an important aspect of theprocess. Other impurities such as arsenic, antimony, etc., are alsoremoved with iron by co-adsorption or precipitation mechanisms.

It has been found that some concentrates, however, exhibit so muchsulphur oxidation (acid generation) that the acid-consuming capacity ofpressure oxidation is exceeded, and some iron is leached into solution,even under process Mode B conditions. It is a target of the process toproduce a low-iron liquor, typically with 0.05 g/L Fe. Some concentrateswhich have been tested have produced pressure oxidation liquors with 1.0to 12.0 g/L Fe. Similarly the pH of the pressure oxidation liquor isnormally targeted to be in the range 2.0 to 3.5, corresponding to lessthan 1 g/L free acid, but concentrates tested have produced pressureoxidation liquors with pH in the range 1.2-2.0, corresponding to 1 to 15g/L free acid.

Accordingly, a further embodiment of the process has been developed,termed “process Mode C” for the treatment of the above concentrates,termed “Mode C” concentrates. Process Mode C will now be describedbelow.

Process Mode C

The Mode C concentrates that exhibit a strong tendency towards sulphuroxidation and hence acid generation are those with a high S:Cu ratio, ormore generally S:M ratio, where M=base metals, such as Cu, Zn, Ni, Co,Pb, etc., but not including Fe, which does not consume acid.

Nickel or nickel/copper concentrates may often by Mode C, because theyare frequently low-grade, with S:M ratio often about 2:1 or higher. Somecopper or copper/gold concentrates are also Mode C, if they are lowgrade because of high pyrite content. Some copper/zinc concentrates havealso been found to be high in pyrite and hence of Mode C type as well.

In general there is a correlation between pyrite (FeS₂) content and thetendency toward Mode C type behaviour. However, there are alsoexceptions to this trend, as not all pyrites react in the same way. Somepyrites oxidize sulphur more readily than others. In contrast,pyrrhotite (Fe₇S₈) or the related iron-zinc mineral sphalerite, (Zn,Fe)S, appear to result in much less sulphur oxidation, and thus exhibitProcess Mode A behaviour.

Process Mode C is essentially a special case of Process Mode B, with twokey features.

First, all the raffinate 63 (FIG. 2) is neutralized, before returningthis stream to the pressure 12 oxidation, (i.e. none is bypassed).

Secondly, the pressure oxidation slurry (before filtration of leachresidue) is subjected to an extra neutralization, the pressure oxidationneutralization, to neutralize excess acid and precipitate any Fe insolution at this time. The pressure oxidation neutralization is done ashot as practical, once the slurry has been discharged from theautoclave. The most convenient opportunity is in the conditioning tankafter flash let-down to atmospheric pressure, when the slurry is at ornear the boiling point of the solution, i.e. about 100° C.

Limerock is used for this purpose, to neutralize any surplus acid in thepressure oxidation slurry and thus bring the pH up to about 3.Simultaneously, any dissolved Fe present in the Fe³⁺ state will beprecipitated, along with any As or Sb that may be present.

The principal products of these reactions are precipitated gypsum andiron hydroxides or basic salts. Since the pressure oxidationneutralization is done before filtration, these solids are mixed in withthe leach residues already present in the pressure oxidation slurry,containing mostly elemental sulphur, hematite, unreacted sulphides(pyrite), and any gangue minerals (quartz, feldspars, etc., which arelargely unchanged by pressure oxidation). This mixing is advantageous asno additional filtration step is required, and the other solids aid inthe filtration of the pressure oxidation neutralization products, whichmight otherwise tend to filter poorly.

The resultant slurry, now at pH 3 is filtered and the filter cakecarefully washed, as always, to remove entrained liquor (Cu, Cl) as muchas practical. The filter cake proceeds to atmospheric leaching where anyprecipitated copper is leached as usual at about pH 1.5-1.8, and theresultant washed thoroughly in a CCD circuit. The filtrate 29 from thepressure oxidation filtration is treated as in Process Mode B for Curemoval by the solvent extraction stage 50, producing a raffinate 63that then goes to neutralization 76 as before, and then recycled back tothe pressure oxidation 12, but without the raffinate split 70, asindicated above. Thus the pressure oxidation cycle is completed.

The important aspects of the process according to the invention can besummarized as follows:

(i) oxidize completely all base metals contained in sulphideconcentrates, e.g. copper, nickel, zinc and cobalt, as well as iron; and

(ii) minimize the oxidation of sulphur to sulphate and maximize theproduction of elemental sulphur; and

(iii) precipitate the metals oxidized during pressure oxidation as thebasic salt, e.g. basic copper sulphate; or

(iv) solubilize the metals oxidized during pressure oxidation, as thesulphate compound, e.g. zinc sulphate or nickel sulphate.

Although the pressure oxidation is chloride catalyzed, it does not use astrong chloride solution, e.g., only about 12 g/L is needed which willsupport about 11 g/L Cu or Zn as the respective chloride salt. If ahigher concentration of metals is needed or produced, it is as thesulphate salt. Thus, the pressure oxidation solutions are generallymixtures of the sulphate and chloride salts, not pure chlorides.

The process according to the invention can be used to processconcentrates containing nickel alone or in combination with copper orcobalt. Similarly, copperzinc concentrates can be processed. This isachieved by the correct use of sulphate or sulphuric acid duringpressure oxidation in the presence of a halogen, such as chloride.Insufficient acid or sulphate increases sulphur oxidation, which isundesirable, as well as reduces metal oxidation and hence metalrecovery. Excess acid solubilizes iron from the pressure oxidationslurry and causes unnecessary expense in cost of acid and neutralizingagent.

Copper-Nickel Concentrates

The process flowsheet is shown as FIG. 5. It is intended forconcentrates containing 3-25% Cu and 3-10% Ni, with Cu predominant.Generally cobalt is present at a Ni:Co ratio of between 10:1 and 30:1,which corresponds to about 0.1 to 0.8% Co in concentrate.

The process essentially is variation of Process Mode B above, where Cureports primarily to the liquor during pressure oxidation, rather thanto the solid product. Acid must be supplied to pressure oxidation toenable both the Ni and Cu to solubilize primarily as sulphate. Typicallyabout 20-30 g/L acid as H₂SO₄ is added to pressure oxidation feedsolution. Chloride addition to pressure oxidation is sufficient tomaintain 12 g/L Cl, same as for Cu concentrates. Conditions oftemperature, pressure, etc., are also similar as for Cu concentrates. Cosolubilizes along with Ni.

The pressure oxidation liquor is treated first for Cu solventextraction, to remove essentially of the Cu, and then Ni is precipitatedas Basic Nickel sulphate, after first reheating to 85 to 90° C., usinglimestone. Co is precipitated along with Ni as a basic cobalt salt.

The precipitated basic nickel/cobalt sulphate is then leached in anammoniacal solution recycled from Ni solvent extraction. The resultantNi/Co leach liquor is then treated first for Co removal by solventextraction using a reagent specific for Co such as Cyanex 272, aproprietary phosphinic acid from Cyanamid Inc. The Co raffinate is thentreated for Ni recovery by another solvent extraction reagent, LIX 84, aproprietary hydroxy-oxime from Henkel Corp.

Finally, the Ni raffinate is recycled to the Ni/Co leach. There is ableed of this raffinate which is treated to recover Ammonium Sulphatewhich otherwise would build up in the circuit. This is due to theintroduction of sulphate ions in the basic nickel sulphate filter cake.Ammonia must be added to make up for the loss of ammonia in the ammoniumsulphate.

Nickel-Copper Concentrates

Nickel-copper concentrates have Ni as the predominant element and willcontain about 8-25% Ni and 3-10% Cu. The process flowsheet is shown inFIG. 6. Conditions in pressure oxidation are essentially the same as forCopper-nickel concentrates. The difference from FIG. 3 lies in thetreatment of the pressure oxidation slurry.

These concentrates generally behave as in Process Mode A, in which Cureports primarily to the solid phase after pressure oxidation. This isaccomplished by addition of limerock to the pressure oxidation slurry toraise the pH to about pH4, before the slurry is filtered. This has theeffect of neutralizing excess acid in pressure oxidation liquor;precipitating any Fe; and precipitating any Cu.

The neutralized slurry is filtered and the filter cake sent toatmospheric leach, labelled a “copper leach” which in turn produces aleach liquor for extraction by Cu solvent extraction.

The neutralized solution is treated for Ni/Co recovery by precipitationand solvent extraction as for copper-nickel concentrates.

Nickel Laterite Ores

Nickel laterites do not concentrate by flotation as sulphides do andtherefore have to be treated as a whole ore. Typically they contain1.5-3.0% Ni, and 0.1-0.3% Co, with negligible Cu. An important featureis Mg content which can be up to 20% Mg, as well as substantial Fecontent. The flowsheet is shown in FIG. 7.

The process is similar to that used for Nickel-copper sulphideconcentrates, except that the absence of Cu means that the leachresidue, after neutralization and filtration can be discarded as it hasnegligible metal values in Cu. There are also important differences inthe conditions used in pressure oxidation: Temperature and pressure aremuch higher at 225° C./450 psig O₂, and much higher acidity at 100 to200 g/L free acid in feed liquor. Chloride content stays the same atabout 12 g/L Cl. Chloride in leach liquor may be supplied as MgCl₂ orHCl.

The other main difference is the need for a Mg removal step. Mg leachesalmost quantitatively into solution during pressure oxidation, resultingin typically 40 g/L Mg per pass. This can be removedevaporation/crystallization for example as MgSO₄.

Copper-Zinc Concentrates

Copper-Zinc concentrates with 20-25% Cu and 1-10% Zn are treated byProcess Mode B type flowsheet, as shown in FIG. 8.

It has been found that excellent extraction of Zn in pressure oxidationcan be achieved so long as enough acid is added to the feed solutionthat the final pH of the slurry is below about pH 2. Otherwise, theconditions are similar as for Cu concentrates, i.e., 150° C. 200 psigO₂, 12 g/L Cl.

In Process Mode B flowsheets, the Cu is primarily solubilized during PO,and must be extracted by solvent extraction 50 This solvent extractionis operated in conjunction with Cu solvent extraction that extracts Cufrom the leach liquor coming from atmospheric leach (“Cu Leach”).

Zinc, having been solubilized during pressure oxidation along with Cu,is precipitated from the Cu raffinate as Basic Zinc sulphate bylimestone at pH 6 and at 85°-90° C. The Zn ppt is then leached by acidraffinate returning from Zn solvent extraction circuit. The Zn leachliquor is then extracted by Zn solvent extraction producing a loaded Znorganic (DEHPA). This organic stream must be carefully purified of Cu,Co, Cd, Cl, etc., before stripping with spent acid from electrowinning.The purification is done by scrubbing the loaded organic using ZnSO₄aqueous solution.

While only preferred embodiments of the invention have been describedherein in detail, the invention is not limited thereby and modificationscan be made within the scope of the attached claims.

What is claimed is:
 1. A process for the extraction of a at least onenon-cuprous metal selected from the group consisting of nickel andcobalt from a metal ore or concentrate containing the non-cuprous metal,comprising the step steps of subjecting the metal ore or concentratecontaining the non-cuprous metal selected from the group consisting ofnickel and cobalt to pressure oxidation in the presence of oxygen and anacidic solution containing halogen ions and a source of bisulphate orsulphate ions to form a product solution of said non-cuprous metal,wherein said source of bisulphate or sulphate ions is at least onemember selected from the group consisting of sulphuric acid and a metalsulphate which that hydrolyzes in said acidic solution, recovering thenon-cuprous metal from the product solution, and recycling a portion ofthe product solution remaining after the recovering step to the pressureoxidation step.
 2. The process according to claim 1, wherein saidhalogen is ions are selected from chlorine and bromine ions.
 3. Theprocess according to claim 1, wherein the non-cuprous metal is zinc andfurther comprising the steps of precipitating zinc from said solution ofnon-cuprous metal in the form of a basic zinc sulphate; separating thebasic zinc sulphate from the remainder of said solution; and leachingzinc from said basic zinc sulphate to produce a solution of zinc ions.4. The process according to claim 3, further comprising the steps ofextracting zinc from said solution of zinc ions by means of an organicextractant; and stripping zinc from said extractant to produce aconcentrated solution of zinc ions for electrowinning.
 5. The processaccording to claim 3, wherein said solution of non-cuprous metal alsocontains copper ions and further comprising the step of removing saidcopper ions from solution prior to said precipitation of zinc as thebasic zinc sulphate.
 6. The process according to claim 5, wherein thecopper ions are removed by solvent extraction.
 7. The process accordingto claim 5, wherein the copper ions are removed by selectiveprecipitation of copper.
 8. The process according to claim 1, whereinthe non-cuprous metal is nickel and further comprising the steps ofprecipitating nickel from said solution of non-cuprous metal in the formof a basic nickel sulphate; separating the basic nickel sulphate fromthe remainder of said solution; and leaching nickel from said basicnickel sulphate to produce a solution of nickel ions.
 9. The processaccording to claim 8, wherein said solution of non-cuprous metal alsocontains copper ions and further comprising the step of removing saidcopper ions from solution prior to said precipitation of nickel as thebasic nickel sulphate.
 10. The process according to claim 9, wherein thecopper ions are removed by selective precipitation of copper.
 11. Theprocess according to claim 1, wherein the non-cuprous metal is cobaltand further comprising the steps of precipitating cobalt from saidsolution of non-cuprous metal in the form of a basic cobalt sulphate;separating the basic cobalt sulphate from the remainder of saidsolution; and leaching cobalt from said basic cobalt sulphate to producea solution of cobalt ions.
 12. The process according to claim 11,wherein said solution of non-cuprous metal also contains copper ions andfurther comprising the step of removing said copper ions from solutionprior to said precipitation of cobalt as the basic cobalt sulphate. 13.The process according to claim 12, wherein the copper ions are removedby selective precipitation of copper.
 14. The process according to claim11, wherein the halogen ion concentration is about 12 g/L.
 15. Theprocess according to claim 1, wherein the non-cuprous metal is a mixtureof nickel and cobalt and further comprising the steps of precipitatingnickel and cobalt from said solution of non-cuprous metal in the form ofbasic nickel and cobalt salts; separating the basic nickel and cobaltsalts from the remainder of said solution; and leaching nickel andcobalt from said basic nickel and cobalt sulphates to produce a solutionof nickel and cobalt ions.
 16. The process according to claim 15 furthercomprising the step of separating said nickel and cobalt ions byselective solvent extraction to produce separate nickel and cobaltsolutions for electrowinning.
 17. The process according to claim 1,wherein the halogen concentration is in the range of about 8 g/L toabout 20 g/L.
 18. The process according to claim 1, wherein said sourceof bisulphate or sulphate ions is generated by the metal ore orconcentrate during said pressure oxidation.
 19. The process according toclaim 2, wherein the halogen ions are chloride ions and theconcentration of chloride is in the range of about 8 g/L to about 20g/L.
 20. The process according to claim 19, wherein the chlorideconcentration is about 12 g/L.
 21. The process according to claim 1,wherein the pressure oxidation is carried out at a temperature betweenabout 130° C. and 150° C.
 22. The process according to claim 1, whereinthe pressure oxidation is carried out under a total oxygen and steampressure of about 690 kPa to about 1380 kPa.
 23. The process accordingto claim 1, wherein the non-cuprous metal is nickel, and the recoveringof the nickel comprises precipitating the nickel from the productsolution to form a nickel precipitate and leaching nickel from thenickel precipitate to produce a nickel solution.
 24. The processaccording to claim 23, wherein the product solution also contains copperand the process further comprises removing the copper from the productsolution prior to the precipitating of the nickel.
 25. The processaccording to claim 23, wherein the metal ore or concentrate alsocontains copper that is leached into the product solution during thepressure oxidation and the process further comprises removing the copperfrom the product solution prior to the precipitating of the nickel. 26.The process according to claim 25, wherein the copper is removed bysolvent extraction.
 27. The process according to claim 25, wherein thecopper is removed by precipitation.
 28. The process according to claim1, wherein the non-cuprous metal is cobalt, and the recovering of thecobalt comprises precipitating the cobalt from the product solution toform a cobalt precipitate and leaching cobalt from the cobaltprecipitate to produce a cobalt solution.
 29. The process according toclaim 28, wherein the product solution also contains copper and theprocess further comprises removing the copper from the product solutionprior to the precipitating of the cobalt.
 30. The process according toclaim 28, wherein the metal ore or concentrate also contains copperwhich is leached into the product solution during the pressure oxidationand the process further comprises removing the copper from the productsolution prior to the precipitating of the cobalt.
 31. The processaccording to claim 30, wherein the copper is removed by solventextraction.
 32. The process according to claim 1, wherein the at leastone non-cuprous metal is both nickel and cobalt, and the recovering ofthe nickel and cobalt comprises precipitating the nickel and cobalt fromthe product solution to form a nickel and cobalt precipitate andleaching nickel and cobalt from the nickel and cobalt precipitate toproduce a nickel and cobalt solution.
 33. The process according to claim32, wherein the product solution also contains copper and the processfurther comprises removing the copper from the product solution prior tothe precipitating of the nickel and cobalt.
 34. The process according toclaim 32, wherein the metal ore or concentrate also contains copper thatis leached into the product solution during the pressure oxidation andthe process further comprises removing the copper from the productsolution prior to the precipitating of the nickel and cobalt.
 35. Theprocess according to claim 34, wherein the copper is removed by solventextraction.
 36. The process according to claim 34, wherein the copper isremoved by precipitation.
 37. The process according to claim 32, whereinthe process further comprises separating the nickel and cobalt from thenickel and cobalt solution by selective solvent extraction to produceseparate nickel and cobalt solutions for electrowinning.
 38. A processfor the extraction of a non-cuprous metal from a metal ore orconcentrate containing the non-cuprous metal, comprising the steps ofsubjecting the metal ore or concentrate containing the non-cuprous metalto pressure oxidation at a temperature of between about 130° C. and 150°C. in the presence of oxygen and an acidic solution containing halogenions and a source of bisulphate or sulphate ions to form a productsolution of said non-cuprous metal, wherein said source of bisulphate orsulphate ions is at least one member selected from the group consistingof sulphuric acid and a metal sulphate that hydrolyzes in said acidicsolution, recovering the non-cuprous metal from the product solution,and recycling a portion of the product solution remaining after therecovering step to the pressure oxidation step.
 39. A process for theextraction of at least one non-cuprous metal selected from the groupconsisting of nickel and cobalt from a metal ore or concentratecontaining the non-cuprous metal, comprising the steps of subjecting themetal ore or concentrate containing the non-cuprous metal to pressureoxidation at a temperature of between about 130° C. and 150° C. in thepresence of oxygen and an acidic solution containing halogen ions and asource of bisulphate or sulphate ions to form a product solution of saidnon-cuprous metal, wherein said source of bisulphate or sulphate ions isat least one member selected from the group consisting of sulphuric acidand a metal sulphate that hydrolyzes in said acidic solution, andrecovering the non-cuprous metal from the product solution.